The flotation of bastnaesite, as a major mineral source of rare earth elements, attracting much attention in the mineral processing field, is challenging owing to the natural flotability of calcium-bearing minerals. To promote the application of flotation, we systematically investigated the flotation behavior of bastnaesite, barite, and calcite, with salicylhydroxamic acid (SHA) as the collector through micro-flotation experiments, zeta-potential measurements, Fourier transform infrared (FT-IR) analyses, X-ray photoelectron spectroscopy (XPS) analyses, and solution chemistry analyses. Micro-flotation experiments confirm that the flotability of bastnaesite is high at pH 6.5–8.5, while calcite floats at pH 8.0–9.5, and barite has little flotation response. The results of FT-IR, XPS, and zeta-potential measurements indicate that there is chemical adsorption of SHA on the bastnaesite surface, and physical adsorption also occurs. However, as for barite and calcite, there is only physical adsorption of SHA on the surfaces. The solution chemistry results show that SHA anions can interact with RE³⁺, REOH²⁺, and RE(OH)₂⁺ on bastnaesite surfaces in aqueous suspensions, resulting in bastnaesite flotation.

During the flotation separation process of bastnaesite, it is difficult to separate bastnaesite from fluorite effectively. In this present study, sodium silicate (SS) can effectively improve the flotation separation effect of bastnaesite and fluorite in salicylhydroxamic acid (SHA) systemasa. Through relevant analyses, such as Zeta potential measurements, adsorption capacity tests, Fourier transform infrared (FTIR) spectroscopic analyses and X-ray photoelectron spectroscopy (XPS) tests, the selective suppressor of SS on fluorite was proven. At pH 10, the single mineral flotation results show that with the increase of SS dosage, the flotation recovery of fluorite rapidly decreases from 61.5% to 35.31%, while the flotation rate of bastnaesite is still high (recovery is 80.02%). Then, the experiment of artificial mixed ore proved that the flotation separation of fluorite and bastnaesite was effective under the appropriate dosage of inhibitor. The results of potentiodynamic measurement and an adsorption capacity test showed that the $\mathrm{SiO}{\left(\mathrm{OH}\right)}_3^{-}$ structure of SS more easily reacted with fluorite, which further prevented the adsorption of SHA on the fluorite surface. FTIR test results and XPS analysis further showed that SS had a strong binding effect with the Ca site on the fluorite surface, but a weak binding effect with the Ce site on the bastnaesite surface. Consequently, SS can be used as an effective inhibitor in the flotation separation of fluorite and bastnaesite.

In this work, selective flotation separation of microfine apatite and chlorite was achieved by using sodium oleate (NaOL) as a collector with a low dosage of acidified sodium silicate (ASS) as a depressant. The optimum ratio of sodium silicate to sulfuric acid for ASS was 5:3, and a good separation effect was also achieved in the mixed ore system. Compared to the deionized water system, the ions of Na⁺, Ca²⁺ and Mg²⁺ in seawater adsorbed on the surfaces of apatite and chlorite, which made the zeta potential of the two minerals shift positively. This presented a challenge to the selection of reagents for mineral separation. The addition of ASS changed the pH value of the pulp from weak alkalinity caused by seawater to weak acidity, which allowed the metal ions adsorbed on the mineral surface to desorb. Meanwhile, ASS can selectively adsorb on the desorbed chlorite surface in the form of Si(OH)₄, which hindered the action of NaOL, leading to the depression of chlorite. NaOL adsorbed well on the desorbed surface of apatite and increased the apatite particle size from 27 μm to 229 μm, with a hydrophobic agglomeration effect, thus enhancing the flotation of microfine apatite.

In this study, 2-hydroxy-3-naphthalene hydroxamic acid (NHA) was used as a collector in microflotation experiments. By comparing the flotation performance of NHA with that of sodium oleate (NaOL) and salicylhydroxamic acid (SHA) commonly used in industry, the performance of NHA in the flotation separation of bastnaesite and calcite was studied. Potentiodynamic (zeta) measurements, Fourier transform infrared spectroscopy (FTIR) measurements and X-ray photoelectron spectroscopy (XPS) measurements were used to reveal the interaction mechanism of bastnaesite with NHA. The results of the pure mineral microflotation test showed that when the pH value was 9 and the dosage of NHA was 4.0 × 10⁻⁴ mol/L, the recovery of bastnaesite reached 93.5%. A concentrate with a bastnaesite grade of 87.08% was obtained from the flotation test of artificially mixed ore, and the recovery of bastnaesite was 90.83%. The zeta measurements and FTIR experiments showed that NHA adsorbed onto the surface of bastnaesite, which changed the surface state of bastnaesite. XPS testing showed that NHA chemically adsorbed onto the bastnaesite, and Ce³⁺ formed a chelated structure with $-\mathrm{C}\left(\mathrm{OH}\right)=\mathrm{NO}-$. The hydrophobicity of the surface of bastnaesite was improved, which made it easier for bastnaesite to adhere to the slurry surface by bubbles. At present, most of the domestic methods for recovering fine-grained bastnaesite use flotation recovery. Calcite and other gangue minerals inevitably enter into the concentrate resulting in low-grade REE concentrate.

To address the marked differences in the host phases of phosphorus and vanadium, severe ore sliming, and the difficulty of synergistic recovery from the low-grade phosphorus–vanadium-associated ore in Mazharweina, Xinjiang, process mineralogical analysis, phosphorus flotation condition tests, and acid leaching-solvent extraction tests were conducted to investigate phosphorus and vanadium recovery from this ore. The results showed that the P₂O₅ and V₂O₅ grades of the raw ore were 16.71% and 0.798%, respectively. Phosphorus occurred mainly in collophane, whereas vanadium did not form independent minerals and was mainly dispersed in carriers such as mica, argillaceous silicates, and colloidal limonite. P₂O₅ and V₂O₅ showed no obvious natural enrichment in any particle-size fraction, and effective separation of P and V could not be achieved by particle-size classification. Therefore, a pre-desliming-direct phosphorus flotation process was adopted. Under the conditions of a grinding fineness of 80% passing −74 μm and an EM-C2 dosage of 1400 g/t, closed-circuit flotation comprising one roughing, one scavenging, and three cleaning stages produced a phosphate concentrate with a P₂O₅ grade of 28.32% and a recovery of 59.50%, while the P₂O₅ grade of the tailings decreased to 1.52%. Vanadium was mainly enriched in the slime and dephosphorization tailings. At a liquid-to-solid ratio of 8:1, an H₂SO₄ mass concentration of 250 g/L, and a leaching time of 8 h, the V₂O₅ leaching rate from the slime was 77.83%. After N235 extraction, Na₂CO₃ stripping, vanadium precipitation, and calcination of the leachate, a vanadium product with a V₂O₅ purity of 98.42% was obtained, with an overall recovery from the raw ore of 63.83%. The results demonstrate the feasibility of combining phosphorus flotation with hydrometallurgical treatment of the slime and dephosphorization tailings for the comprehensive recovery of phosphorus and vanadium from this complex phosphorus–vanadium ore.